Flotation of heavy metal oxides



United States Patent h 7 2,861,687 FLOTATION OF HEAVY METAL OXIDES Joseph Robert Lord, Los Angeles, Calif., assignor to Southwestern Engineering Company, a corporation of California No Drawing. Application September 9, 1955 Serial No. 533,524 32 Claims. (Cl. 209-167) My invention relates to a process for the concentrating of the oxide ores of heavy metals by flotation. In particular, my process is applicable to the ores of heavy metals such as titanium, arsenic, antimony, bismuth, van'adium, tantalum, tungsten, molybdenum, iron, manganese,

chromium, tin, uranium, zirconium and the like. While the process is readily adaptable to both slimin'g and nonsliming ores it is particularly advantageous in the case of the former since no de-sliming step. is required, as has been heretofore necessary in conventional soap flotation processes applied to such ores.

The object of my invention is to concentrate such an ore by a method which is eflicient, simple and economical from the standpoint of grade of concentrate and recovcry, and also from the standpoint of the cost of the re agents consumed. It is also an object of my invention to produce from a heavy metallic oxide ore a concentrate which is either of marketable grade or may be made of marketable grade by nodulizing or sintering.

The aforementioned and other objects are accomplished in accordance with my method by incorporating a novel combination of reagents into a pulp containing the ore in powdered form, the novel combination of reagents being an organic acid collector such as fatty acids having :at least 12 carbon atoms and their water soluble soaps,

or naphthenic acids and their water soluble soaps; a waiter soluble salt of a heavy metal; and a hydrocarbon oil.

The following examples set forth in detail specific embodiments of my invention, and are to be considered not by way of limitation. For the various examples, the statement regarding percentage by weight of manganese (or other metal) contained in the sample to be tested is based on assay of the sample before testing, whereas the statement regarding percentage of recovery is based on calculated percentage of manganese (or other metal) contained in that portion of the general sample which has been used for the particular test. This method of expressing recovery, based on calculated head rather than assayed head, is common practice to compensate for slight variations in mineral content of the individual portions of the general sample used for the various tests. Where I have employed fuel oil as the hydrocarbon oil, I refer to a petroleum derivative having the following specification, although others may of course be used:

Market name: Automotive Diesel Fuel. Gravity: 31 to 33 A. P. I.

Viscosity 122 F.-42 seconds. A lflphur00.50%. 1 P

2,851,687 Patented Nov. 25, 1958 2 EXAMPLE I The ore employed in this test was manganese ore obtained from the Three Kids Mine, Las Vegas Wash, Nevada. wad with minor amounts of pailomelane and man'ganite, and the gangue is chiefly quartz, gypsum, decomposed tufl, mica and various other silicatea. The ore is a sliming oxide ore and the sample used contained 25.12% by weight of manganese.

A sample of the ore which had been crushed to -10 mesh was formed into a slurry containing 40% by weight of solids with softened tap water by grinding in a ball mill for 5 minutes at ordinary temperature, the ore then having the following sieve analysis: not passing a mesh sieve, 0.75% by weight; passing a 65 mesh sieve but not passing a mesh sieve, 3.50%; passing a 100 mesh sieve but not passing a mesh sieve, 7.50%; pass! ing a 150 mesh sieve but not passing a 200 mesh sieve, 9.50%; passing a 200 mesh sieve but not passing a 325 mesh sieve, 15.50%; and passing a 325 mesh sieve, 63.25%; whereby 78.75% was 200 mesh.

After grinding, a quantity of the slurry containing 200 grams of the powdered ore was placed in a 500 gram laboratory Fagergren flotation machine, and a further quantity of the softened tap water was added to form a pulp containing 9% by weight of the powdered ore. The following reagents were then added separately in the following amounts (pounds per ton of ore), respectively: sodium oleate, 25.5; domestic fuel oil, 56.0; Oronite S (a proprietary product, being the sodium salt of the sulfonation product of the reaction product of tetraisopropylene and a lower homologue of benzene), 4.0; and Mn'SO -2H O, 20.0. The pulp wasthen conditioned for a period of 3 minutes at ordinary temperatures with air valve of test cell closed.

At the end of the conditioning period, air was admitted for 3 minutes to produce a rougher concentrate and tailings. The tailings amounted to 26.1%, based upon the weight of the ore sample taken, and contained 6.02% of manganese.

The trailings were removed from the apparatus and discarded, and the rougher concentrate was then subjected to further flotation in the cell using the softened water as make-up as needed. Sodium silicate in the amount of 2.5 pounds per ton, based upon the weight of the original ore sample taken, was placed in the cell. At this point, the material in the cell had a pH of 8.0. The cell was then operated for a period of 3 minutes to provide a concentrate amounting to 61.9% based upon the weight of the ore sample taken. A cleaner tailing (middling No. l) was produced amounting to 12.0% based upon the weight of the ore sample taken.

This second concentrate was further subjected to flotation in the cell for a period of 3 minutes and produced a concentrate amounting to 57.6% based upon the weight of the ore sample taken. A cleaner tailing (middling No. 2) was'produced amounting to 4.3% based upon the weight of the ore sample taken. This third concentrate was further subjected to flotation in the cell for a period of 3 minutes and produced a final concentrateamounting to 54.3% based upon the weight of the In this ore, most of the manganese occurs as 3 combined first, second an'dthird middlings, and 6.2% in the tailings.

EXAMPLE II To illustrate the application of my process to a nonsliming manganese oxide ore, several experiments were performed on samples taken from ore deposits in the Artillery Peaks Mountains, Mohave County, Arizona, located some fifty-five miles south of Yucca. The manganese minerals in these deposits have been identified as psi lomelane and pyrolusite, with minor amounts of braunite and manganite intimately associated with quartz, chalcedony, calcite and barite gangue. Owing to the intimate association of the manganese and a portion of the gangue, and owing to the possibility that the impurities are actually included in the manganese mineral particles, it is recognized that only medium-grade concen- {rates (28-33%rnanganese content) can be produced from such deposits with reasonably high recovery.

pair of experiments was performed on identical samples which assayed 13.70% by weight of manganese. The two samples were similarly crushed and in general theprocess followed the same procedure as employed in Example I except that all reagents were introduced to the ball mill at the start of grinding and the pulp was diluted to 17% by weight of solids of which less than 1% remained on a 65-mesh screen, grinding time being approximately 18 minutes. l A

One sample was used without manganous sulfate and the other employed 5 pounds of such material per tonof ore. The other reagents employed were the same as in Example '1 except that 2.25 pounds of Swift hard soap per ton of ore was used in place of sodium oleate. (Swift hard soap is a proprietary product, made from an acidulated soap stock from mixed vegetable oils comprising principally cotton seed oil.) Approximately half the proportion of wetting agent, fuel oil and sodium silicate of Example I were employed.

Conditioning was carried out for 17 minutes by grinding in the ball mill, and in producing the rougher concentrate the cell was operated for a period of 6 minutes. In producing the first middling, second middling third middling and final concentrate the cell was operated for periods of 4 minutes, 3 minutes and 3 minutes, respectively.

In the experiment omitting the manganous sulfate, the concentrate amounted to 18.7%, based upon the weight of the ore taken, and analyzed 32.65% by weight ofmanganese and 11.75% of silica. The first middling and the combined second and third middlings amounted to 16.0% and 12.7 respectively, and analyzed 12.49% and 22.28% manganese respectively. Thetailings amounted to 52.6% and analyzed 5.54% of manganese. V v

Thus, only 44.2% of the manganese was recovered in the concentrate while 34.95% and'2 l% were in the combined middlings and tailing, respectively. A

On the other hand, in the experiment employingmanganous sulfate in which an identical procedure was followed, the concentrate amounted to 31.9%, based upon the weight of the ore, and analyzed 31.32% manganese and 18.08% silica. The first middling, and the combined second and third middlings amounted to 11.5% and 10.1%, respectively, and analyzed 6.61% and 13.11% manganese. The tailing amounted to 46.4% and analyzed 3.96%.

"Thus,'71.'8% of the manganese present in the ore was recovered in the final concentrate with the loss of only 15.0% and 13.2% manganese inthe combined middlings and tailing, respectively.

oxide ore slime fraction recovered from tailing pond fines in an iron ore washery (Hawkins Washery operated 4 by Cleveland Cliffs Iron Co.) assaying 11.52% iron was treated.

A first 3-minute conditioning period was employed in which a pulp, diluted to 10% by weight of solids, was treated with 2.69 pounds of sodium oleate per ton of ore, 11.3 pounds of fuel oil per tone of ore, 0.9 pound per ton of Oronite S and 4.48 pounds per ton of dried, powdered ferrous sulfate. A second conditioning period of 3 minutes followed in order to deactivate and defio-cculate the silica in the ore. For this purpose 6.27 pounds per ton of sodium fiuosilicate and 5.20 pounds per ton of sulfuric acid were used. The cell was then operated for 3 /2 minutes to produce a rougher concentrate which was refloated for 3 minutes to produce a final concentrate.

The final concentrate amounted to 19.2% based upon the weight of the ore sample and analyzed 46.20% iron. The single middling amounted to 9.8% and analyzed 8.12% iron, and the tailing amounted to 71.0% based upon the weight of the ore sample and analyzed 3.42%

iron.

Thus, the concentrate contained 73.4% of the iron present in the ore whereas the middling and tailing contained respectively 6.6% and 20.0%.

The following examples illustrate the effect of employ- EXAMPLE IV In this example, the procedure of Example I was substantially repeated. In place of the MnSO -2H O, however, CuSO -5H O in the same molar amount was used. The conditioning time was 3 minutes, but in producing the rougher concentrate the cell was operated for only 2% minutes. Also, in producing the first middling, second middling, third middling and final concentrate, the cell was operated for periods of 2 /4, 2% and 2% minutes, respectively.

The concentrate amounted to 58.5%, based upon the weight of the ore taken. The first, second and third middlings amounted to 10.1%, 4.9% and 2.3%, respectively, and combined theyv analyzed 12.53% manganese. The tailings amounted to 24.2% and analyzed 6.19% of manganese. The concentrate analyzed 37.6% manganess and 11.90% silica.

Thus, 85.7% of the manganese present in the ore was present in the concentrate, wher eas 8.5% and 5.8% were present in the combined middlings and tailings, respectively.

EXAMPLE V In this example, the procedure of Example I was substantially repeated. However, in placeof the there was used Pb(NO in the same molar amount.

The conditioning was performed for a period of 3 minutes, whereas in producing the rougher concentrate the cell was operated for 3% minutes. Also, in producing the first middling, second middling, third middling and final concentrate, the cell Was operated for periods of 3 /2 minutes, 3 minutes and 3 minutes, respectively.

The concentrate amounted to 53.8%, based upon the weight of the ore, and analyzed 37.00% manganese. The first, second and third middlings amounted to 14.6%, 4.0% and 2.6%, respectively, and when combined, analyzed 15.10% of manganese. The tailings amounted to 25.0% and analyzed 3.93% of manganese. The concentrate analyzed 11.20% of silica.

Thus, 82.6% of the manganese present in the ore was present in the concentrate, whereas 13.3% and 4.1% were present in the combined middlings and tailings, respectively.

EXAMPLE VI In this example, the procedure of Example I was substantially repeated. Instead of MnSO -2HgO,-"however, the same molar amount of FeSO -7H O was used.

The conditioning period was 3 minutes, whereas in producing the rougher concentrate the cell was operated for 3% minutes. In producing'the first middling, second middling, third middling and final concentrate, the cell was operated for periods of 3% minutes, 3% and 3% minutes, respectively. l

The concentrate amounted to 67.1%, based upon the weight of the ore taken, and analyzed 34.43% of manganese. The first, second and third middlings amounted to. 9.3%, 4.0% and 2.4%, respectively, and combined they analyzed 6.15% of manganese. The tailings amounted to 17.2% and analyzed 4.48% of manganese. The concentrate analyzed 15.30% of silica.

Thus, 93.0% of the manganese present in the ore was present in the concentrate, whereas the amounts present in the combined middlings and tailings were 3.9% and 3.1%, respectively.

EXAMPLE VII In this example, the procedure of Example I was substantially repeated. Instead of MnSO -2H O, however, the same molar amount of Fe(NO -9H O was used.

The conditioning period was 3 minutes, whereas in producing the rougher concentrate the cell was operated for 6 minutes. In producing the first middling, second middling, third middling and final concentrate the cell was operated for periods of 4 minutes, 4 minutes, and 4% minutes, respectively.

The concentrate amounted to 60.4%, based upon the weight of the ore taken, and analyzed 35.14% of manganese and 13.8% of silica. The first, second and third middlings amounted to 9.4%, 3.0% and 2.1%, respectively, and combined they analyzed 9.36% of manganese. The tailings amounted to 25.1% and analyzed 10.00% of manganese.

Thus, 84.6% of the manganese present in the ore was present in the concentrate, whereas the amounts present in the combined middlings and tailings were 5.4% and 10.0%, respectively.

EXAMPLE VHI In this example the procedure of Example I was substantially repeated. Instead of MnSO -2H O, however, the same molar amount of cupric acetate was used. Also, the manganese ore used was of a slightly different grade taken from the same source and assayed 24.89% by weight of manganese.

The conditioning period was 3 minutes, whereas in producing the rougher concentrate the cell was operated for 3 /2 minutes. In producing the first middling, second middling, third middling and final concentrate, the cell was operated for periods of 3% minutes, 3 /2 minutes and 3 /2 minutes, respectively.

The concentrate amounted to 42.8%, based upon the weight of the ore taken and analyzed 39.42% of manganese. The first, second and third middlings amounted to 15.9%, 3.6% and 1.4%, respectively, and combined they analyzed 20.73% of manganese. The tailings amounted to 36.3% and analyzed 9.78% of manganese. The concentrate analyzed 8.50% of silica.

Thus, 68.3% of the manganese present in the ore was present in the concentrate whereas the amounts present in the combined middlings and tailings were 17.4% and 14.3%, respectively.

EXAMPLE IX In this example, the procedure of Example I was substantially repeated. However, in place of the .used in Example I there was used instead the same molar amount of MIICIQ'4H20. The conditioning time was 3 minutes, but in producing the rougher concentrate, first ganese.

ganese. The tailings amounted to 26.6% and analyzed 6.07% of manganese. The concentrate analyzed 9.71% of silica.

Thus, the concentrate contained 85.0% of the manganese present in the ore, whereas the combined middlings and tailings contained, respectively, 8.5% and 6.5%.

In practicing the present process, incorporation in the pulp of a water-soluble salt of a heavy metal is essential. Such incorporation is ordinarily by addition to the pulp of the chosen heavy metal salt, as purchased for example commercially. Such incorporation may be brought about also, in the event that the heavy metal -of the chosen heavy metal salt is the same as the heavy metal of the heavy metal oxide ore undergoing beneficiation, by deriving the water-soluble heavy metal salt directly from the ore by treating with a suitable chemical reagent which will react with the heavy metal oxide ore in such a fashion as to give the desired water-soluble heavy metal salt. For example, when a manganese oxide ore is being floated, addition of sulfurous acid will result in the formation of manganous sulfate. With other ores, for example many iron oxide ores, addition of sulfuric acid may be used to give a water-soluble iron su1-- fate; and similarly with other heavy metals. It will be understood that only a very small part of the oxide ore isso subjected to chemicalreaction; merely enough to. provide a suitable quantity of heavy metal ions.

As stated, the heavy metal salt is essential. Thus, an

experiment was performed in a manner similar to that The concentrate amounted to 11.4%, based upon the weight of the ore sample, and analyzed 18.88% of man- The first middling, second middling and third middling amounted to 17.3%, 13.9% and 9.2%, respectively, and combined they analyzed 26.40% of manganese. The tailings amounted to 48.2% and analyzed 26.00% of manganese.

Thus, 8.5 of the manganese present in the ore was present in the concentrate, 42.0% in the combined middlings and 49.5% in the tailings.

Another experiment was also performed in a manner similar to that set forth in Example I, but using Ca(NO -4H O in place of the MnSO -2H O, but in the same molar amount. The conditioning time used was 3 minutes, and the period of time used in producing the rougher concentrate was 9 minutes. In this experiment efiective flotation did not take place, the eifect being similar to that observed in the experiment just described,

A further experiment was also performed in a man-' ner similar to that set forth in Example I, but using Na SO -10H O in place of MnSO -2H 0 in the same molar amount. In this test, no satisfactory result was obtained, the effect being substantially the same as when Thefollowingsix .examplesillustr'ate the effect of using collectors other than the sodium ,oleate employed in Example I.

EXAMPLE X In this example the procedure of Example I was substantially repeated. Instead of sodium oleate, however, the same molar amountof oleic acid was used. Also the manganese ore used was of a slightly different grade taken from the same source and assayed 24.89% by weight of manganese.

The conditioning period was 3 minutes, whereas in producing the rougher concentrate the cell was operated for 7 minutes. In'producing the first middling, second middling, third middling and final concentrate, the cell was operated for periods of 4% minutes, 2% minutes, and 3 minutes respectively.

The concentrate amounted to 36.2% based upon the weight of the ore taken and analyzed to 38.54% of manganese and 9.30% of silica. The first, second, and third middlings amounted .to 18.5%, 7.3% and 2.8%, respectively, and combined they analyzed 18.25% of manganese. The tailing amounted to 35.2% and analyzed 16.08% of manganese.

Thus, 56.2% of the manganese present in the ore was present in the concentrate whereas the amounts present in the combined middlings and tailing were 21.0% and 22.8% respectively.

EXAMPLE XI In this example the procedure of Example I was substantially repeated. Instead of sodium oleate, however, the same molar amount of oleic acid emulsified with Oronite S was used. Also the manganese ore used was of a slightly different grade taken from the same source and assayed 24.89% by weight of manganese.

The conditioning period was 3 minutes whereas in producing the rougher concentrate the cell was operated for 7 minutes. In producing the first middling, second middling, third middling and final concentrate the cell was operated for periods of 4 minutes, 3% minutes and 3 minutes, respectively.

The concentrate amounted. to 35.0% based upon the weight of the ore taken and analyzed 37.94% of manganese and 10.40% of silica. The first, second, and third middlings amounted to 16.4%, 7.6% and 2.2%, respectively, and combined they analyzed 19.86% of manganese. The tailing amounted to 38.8% and analyzed 16.79% manganese.

Thus, 53.2%. of the manganese present in the ore was present in the concentrate while the amounts present in the combined middlings and tailings were 20.8% and 26.0% respectively.

EXAMPLE XII In this example the procedure of Example I was substantially repeated. Instead of sodium oleate, however, the same weight of cotton seed oil-foots was used. Also the manganese ore used was of a slightly different grade taken from the same source and assayed 24.89% by weight of manganese.

The conditioning period was 3 minutes, whereas in producing the rougher concentrate the cell was operated for 4 minutes. In producing the first middling, second middling, third middling and final concentrate the cell was operated for periods of 3% minutes, 3% minutes and 3 /2'1ninutes, respectively.

The concentrate amounted to 34.0% based upon the weight of the ore taken and analyzed 40.59% of manganese and 6.2% of silica. The first, second, and third middlings amounted to 10.5%, 4.2%, and 3.0%, respectively, and combined they analyzed 21.61% of manganese. The tailing amounted to 48.3% and analyzed 14.89% manganese.

Thus, 51.5% of the manganesezpresentjn the ore was present in the concentrate WhiiQ-thQBIIIOUDISTPI'BSBIII in the combined middlings and tailings were 23.0% and 25.5% respectively.

EXAMPLE XIII In this example the procedure of Example I was substantially repeated. Instead of sodium oleate, however, 23.7 pounds'of fish oil fatty acids per ton of ore were used. Also the manganese ore used was of a slightly different grade taken from the same source and assayed 24.89% by weight of manganese.

The conditioning period was 3 minutes, whereas in producing the rougher concentrate the cell was operated for 7 /2 minutes. In producing the first middling, second middling, third middling and final concentrate the cell was operated for periods of 3 minutes, 3 minutes and 3 minutes, respetcively.

The concentrate amounted to 26.5% based upon the weight of the ore taken and analyzed 39.00% of manganese and 7.40% of silica. The first, second, and third middlings amounted to 17.4%, 5.5% and 2.1%, respectively, and combined they analyzed 22.48% of manganese. The tailing amounted to 48.5% and analyzed 18.83% manganese.

Thus, 41.2% of the manganese present in the ore was present in the concentrate whereas the amounts present in the combined middlings and tailings were 22.4% and 36.4% respectively.

EXAMPLE XIV In this example the procedure of Example I was substantially repeated. Instead of sodium oleate, however, an equal weight of naphthenic acid soap was used.

The conditioning period was 3 minutes whereas in producing the rougher concentrate the cell was operated for 7 minutes. In producing the first middling, second middling, third middling and final concentrate the cell was operated for periods of 7 /2 minutes, 7 minutes and 7 minutes, respectively.

The concentrate amounted to 53.9% based upon the weight of the ore taken and analyzed 35.62% of manganese and 13.40% of silica. The first, second, and third middlings amounted to 11.0%, 6.5%, and 5.1%, respectively, and combined they analyzed 18.40% of manganese. The tailing amounted to 23.5% and analyzed 6.72% manganese.

Thus, 77.0% of the manganese present in the ore was present in the concentrate while the amounts present in the combined middlings and tailings were 16.7% and 6.3% respectively.

EXAMPLE XV In this example the procedure of Example I was substantially repeated. the same weight of fish oil fatty acid soap was used. Also themanganese ore used was of a slightly different grade taken from the same source and assayed 24.89% by weight of manganese.

The conditioning period was. 3 minutes whereas in producing the rougher concentrate the'cell was operated for 3% minutes. In. producing the first middling, second middling, third middling. and finalconcentrate the cell was operated for periods of 4% minutes, 4 /2 minutes and 4%. minutes, respectively.

The concentrate amounted: to 45.1% based upon the weight of the ore taken and'analyzed 39.95% of manganese'and 8.20% of'silica. The first, second, and third middlings amountedto 10.4%, 46% and 3.3%, respectively, and combined they analyzed 20.88% of manganese. The tailing amountedito 3' 6.6% and analyzed 7.59% manganese.

Thus, 73.2% of 'themanganes'e. present in the ore was present in the concentrate while the amounts present in the combined middlings and tailings were 15.5% and 11.3 %v respectively.

The. following two examples illustrate the effectotvarying-pulp; density; on the ore used 5 in Example I.

Instead of sodium oleate, however,

9 EXAMPLE XVI In this example the procedure of Example I was substantially repeated. However, the pulp was diluted to solids, and the manganese ore was of-a slightly different grade taken from the same source and assayed 24.89% manganese.

The conditioning period was 3 minutes whereas in producing the rougher concentrate the cell was operated for 2% minutes. In producing the first middling, second middling, third middling and final concentrate, the cell was operated'for periods of 2%. minutes, 2% minutes and 3 minutes, respectively.

The concentrate amounted to 39.0%, based upon the weight of the ore taken, and analyzed 39.62% of manganese and 8.50% of silica. The first, second, and third middlings amounted to 8.5%, 6.2% and 6.0%, respectively, and combined they analyzed 23.36% of manganese. The tailing amounted to 40.3% and analyzed 9.49% manganese.

Thus, only 64.1% of the manganese present in the ore was present in the concentrate, while the amounts present in the combined middlings and tailings were 20.1% and 15.8%, respectively.

EXAMPLE XVII In this example the procedure of Example I was substantially repeated with the exception that the pulp was I diluted to 25% solids, and the manganese ore was of a slightly different grade taken from the'same source and assayed 24.89% manganese.

The conditioning period was 3 minutes, whereas in producing the rougher concentrate the cell was operated for 4 minutes. In producing the first middling, second middling, third middling and final concentrate the cell in the combined middlings and tailings were 5.0% and 2.4%, respectively.

In practicing the present process, it is essential for best results, that the gangue dispersion be achieved before the flotation is carried out. Thus, an experiment was performed in substantially the same manner as that of Example I, but omitting the wetting agent.

The conditioning time was 3 minutes, whereas in producing the rougher concentrate, first middling, second middling, third middling and final concentrate, the cell was operated for periods of 5 /2 minutes, 5 minutes, 4 minutes and 3 /2 minutes, respectively.

The concentrate amounted to 40.3%, basedupon the weight of the ore, and analyzed 38.88% of manganese. The first middling, second middling and third middling amounted to 13.4%, 6.0% and 5.0%, and combined they analyzed 20.16% of manganese. The tailings amounted to 35.3% and analyzed 12.67% of manganese. The concentrate analyzed 11.20% of silica.

Thus-only 62.6% of the manganese present in the ore was present in the. concentrate, whereas the combined middlings and tailings contained 19.6% and 17.8%, respectively.

The following four examples illustrate the effect of varying the relative proportions of fatty acid collector and heavy metal salt, and also the effect of admixing reagents before they are incorporated into the pulp.

' EXAMPLE XVIII stantially repeated. However, the amount of sodium oleate used was 30.6 pounds per ton of ore.

' The conditioning time was 3 minutes, but in producing the rougher concentrate, first middling, second middling, third middling and final concentrate, the cell was operated for periods of 3 minutes, 2% minutes, 2 /2 minutes and 2 /2 minutes, respectively.

The concentrate amounted to 61.5%, based upon the weight of the ore, and analyzed 36.90% of manganese. The first middling, second middling and third middling amounted to 8.1%, 5.2% and 3.0%, respectively, and combined they analyzed 10.54% of manganese. The tailings amounted to 22.2% and analyzed 4.35 of manganese.

Thus, 89.4% of the manganese present in the ore sample was present in the concentrate, whereas the corresponding figures for the combined middlings and the tailings were 6.8% and 3.8%, respectively.

EXAMPLE XIX In this experiment the procedure of Example I was substantially repeated. However, sodium oleate in the amount of 30.0 pounds per ton of ore was used. No sodium silicate was used for cleaning.

The conditioning time was 3 minutes, but in producing the rougher concentrate, first middling, second middling, third middling and final concentrate, the cell was operated for periods of 2%. minutes, 2 /2 minutes, 2 /4 minutes and 3 minutes, respectively.

The concentrate-amounted to 65 .2%, based upon the weight of the ore, and analyzed 36.30% of manganese. The first, second and third middlings, amounted to 7.0%, 4.3%, and 2.7%, and combined they analyzed 8.33% of manganese. The tailings amounted to 20.8% and analyzed 4.52% of manganese. The concentrate analyzed 15.20% of silica.

Thus, 91.8% of the manganese present in the ore was present in the concentrate, whereas the combined middlings and tailings contained 4.6% and 3.6%, respectively.

EXAMPLE XX The procedure of Example I was again substantially repeated, but the amount of MnSO -2H O used was only 7.9 pounds per ton of ore.

The conditioning time was 3 minutes, but in producing the rougher concentrate, first middling, second middling, third middling and final concentrate, the cell was operated for periods of 2% minutes, 2 minutes, 2 minutes and 2 minutes, respectively. Sodium silicate in the amount of 2.5 pounds per ton was also added in the first and second cleaning steps.

The concentrate amounted to 37.8%, based upon the weight of the ore, and analyzed 38.08% of manganese. The first middling, second middling and third middling amounted to 16.0%, 10.8% and 8.3%, respectively, and combined they analyzed 22.88% of manganese. The tailings amounted to 27.1% and analyzed 9.80% of man ganese. The silica content of the concentrate was 10.10%

Thus, the concentrate contained only 57.4% of the manganese present in the ore, whereas the combined middlings and tailings analyzed 32.0% and 10.6%, re spectively.

EXAMPLE XXI The procedure of Example I was substantially repeated. However, the sodium oleate was used in the amount of 30.0 pounds per ton and the MnSO -2H O was used in The concentrate amounted to 55.0%, based upon the weight of the ore, and analyzed 36.24% of manganese. The first, second and third middlings amounted to 7.2%, 4.4% and 3.6%, respectively, and combined they analyzed 17.36% of manganese. The tailings amounted to 29.8% and analyzed 8.41% of manganese. The silica analysis of the concentrate was 12.80%. g

Thus, 79.5% of the manganese present in the ore was present in the concentrate, the corresponding figures for the combined 'middlings and tailings being 10.5% and 10.0%.

EXAMPLE XXII To illustrate effectiveness of the process on tin oxide ores, a gravity concentration tailing from a Bolivian tin ore, containing tin as cassiterite SnO was treate A sample assaying 0.60% Sn was ground to 68.2% 200 mesh in a water pulp at 50% solids by weight, and was floated at 10% solids by weight, with 10 minutes conditioning and 10 minutes flotation.

During conditioning, the following reagents were introduced in the following amounts (pounds per ton of ore) tall oil soap, 7.2; gas oil No. 1, 44.0; Oronite S, 0.25; and lead acetate, 5.0; and sodium lignin sulfonate, 0.5 was added for dispersion and depression of gangue'.

The concentrate amounted to 44% and the tailing amounted to 56%, based upon weight of the head sample. The concentrate and tailing respectively analyzed 0.90% and 0.50% Sn. 60.5% of the total Sn was in the concentrate, with 39.5% in the tailing.

Another sample of the same tin ore was treated in the same manner, with the same reagents, in approximately the same amount, except lead acetate was omitted and was not replaced by another heavy metal salt.

The concentrate and tailing respectively amounted to 33.15% and 66.85%, based upon weight of the head sample. The concentrate and tailing respectively analyzed 0.74% and 0.45% Sn; and 44.9% of the total Sn was in the concentrate, with 55.1% in the tailing.

Thus, when employing a heavy metal salt the sample was up-graded from 0.60% to 0.90% Sn, but when heavy metal salt was omitted it was only up-graded to 0.74% Sn. When employing heavy metal salt, 60.5% of the total Sn was recovered, but when heavy metal salt was omitted only 44.9% of the total Sn was recovered.

The samples were a gravity concentration tailing low L in tin, so that concentration by any method is necessarily low in tin, by assay; but in terms of recovery (when employing reagents including a heavy metal salt), the Sn concentrate, assaying only 0.90% Sn, represents 60.5% of the total Sn of the ore, thereby illustrating applicability of theprocedure to tin oxide ores.

EXAMPLE XXIII To illustrate the eflectiveness of the process on tungsten oxide o'res,-molybdenite flotation tailing containing tungsten as wolframite (FeO, MnO) W0 was treated. Wolframite is sometimes classified as a tungstate, but it is an oxide ore as termed herein since the tungsten occurs as the oxide.

A sample (lot 3245, test 23) assaying 0.033% W05, and which was 43.9% 200 mesh, Was floatedat 45% solids by weight. A pyrite concentrate was first floated; and the tailing containing wolframite was then floated with 7 minutesconditioning, 9 minutes rougher flotation, 3 minutes conditioning, 4 minutes secondrougher flotation, 2- minutes" conditioning, 7 minutes third rougher flota'tion, '2minutesconditioning, 4 minutes first cleaning of" concentrate, 4 minutes conditioning and 14 minutes second cleaning of concentrate.

Usual reagents were employed for first floating pyrite, in'the following, amounts (pounds per ton of Ore): sulfuri'cacid, 2.70; isopropyl Xanthate, 0.03; and Dupont frother B423, 0.04; I

Certain of the reagentsfor subsequently floatingtungsten were added in therou'gher" and cleaning condition- 12 ings, totalling the following amounts (pounds per ton of ore): Crude tall oil (as soap), 4.41; gas oil No. 1, 12.94; and Emcol No. 5100 (wetting agent), 1.56. Erncol No. 5100 is a proprietary product, being an alkanolamine fatty acid condensate. Ferrous sulfate, 0.77 pound per ton of ore, was added in the first rougher conditioning. Sodium hexametaphosphate, totalling 1.16 pounds per ton of ore, was added in the rougher and cleaning conditionings, for dispersion and depression of gangue.

The pyrite concentrate amounted to 4.4%, the tungsten concentrate amounted to 8.0%, the tungsten cleaner tailings amounted to 13.1% and the tungstenrougher tailing amounted to 74.5%, based on weight of the head sample.

The pyrite concentrate analyzed 0.042% W0 the tungsten concentrate analyzed 0.245% W0 the tungsten cleaner tailings analyzed 0.0098% W0 and the tungsten rougher tailing analyzed 0.011% W0 Thus, a tungsten concentrate was produced, up-graded from 0.033% W0 to 0.245% W0 63.3% of the total W0 was recovered in the tungsten concentrate; with 6.0% in the pyrite concentrate and 30.7% in the combined cleaner and rougher tailings.

Another sample of the same molybdenite flotation tailing (lot 3245, test 24) was treated in the same manner, with the same reagents, in approximately the same amounts, except crude tall oil (4.00 pounds per ton of ore) and gas oil No. 1 (10.27 pounds per ton of ore) were emulsified and added in the rougher and cleaning conditionings, as a soap-oil emulsion.

The pyrite concentrate amounted to 4.3%, the tungsten concentrate amounted to 8.2%, the tungsten cleaner tailings amounted to 10.9% and the tungsten rougher tailing amounted to 76.6%, based upon weight of the head sample. 7

The pyrite concentrate analyzed 0.035% W0 the tungsten concentrate analyzed 0.253% W0 the tungsten cleaner tailings analyzed 0.0133% W0 and the tungsten rougher tailing analyzed 0.01% W0 Thus, a tungsten concentrate was produced, up-graded from 0.033% W0 to 0.253% W0 66.2% of the total W0 was recovered in the tungsten concentrate; with 4.8% in the pyrite concentrate and 29.0% in the combined cleaner and rougher tailings.

In each of the examples of concentration of tungsten; the sample was a molybdenite flotation tailing extremely low in tungsten so that concentration by any method is necessarily low in tungsten, by both weight percentage and assay; but in terms of recovery the W03 concentrates (amounting to weight percentage of only about 8% and assaying only about 0.25% W0 represent 6366% of the total W0 of the ore, thereby illustrating applicability of the procedure to tungsten oxide ores.

EXAMPLE XXIV To illustrate eifectiveness of the process on chromium oxide ores, a sample of sub-grade chrome ore from Yreka, California was treated. The principal chrome mineral identified was picotite (Mg, Fe) (Al, Cr) 0 A sample assaying 36.91% Cr O was ground to 64.95% 200 mesh in a water pulp at 71.5% solids by weight; and the ground ore was screened wet on 200 mesh, resulting in sand and slime fractions which were treated separately by flotation. v

The sand fraction was floated at 23% solids by weight, with. 4 minute conditioning, 4 minute rougher flotation, and 3 minute cleaning of rougher concentrate. The following reagents were added during conditioning, in the following amounts (pounds per ton of ore): Swift hard soap, 3.26; fuel oil, 9.10; Oronite S, 0.72; and ferrous sulfate, 0.36. The pulp was made acid with sulfuric and hydrofluoric acid.

The slime fraction was floated at 9% solids by weight, with 5 /2 minute's conditioning, 6'minute rougher flota- 13 tion, and the concentrate cleaned three times for 4 minutes each. The following reagents were added during conditioning, in the following amounts (pounds per ton of ore): Swift hard soap, 12.08; fuel oil, 22.60; Oronite S, 2.68; and ferrous sulfate, 0.90; and A. C. Reagent No. 610 was added as follows: Cleaner No. 1, 0.80; Cleaner No. 2, 0.09; and Cleaner No. 3, 0.09, for depression and dispersion of gangue. American Cyanamid Reagent No. 610 is a proprietary product which is a lignin sulfonate.

The products from separately floating the sand and slime fractions were combined, with the combined cleaner concentrates amounting to 65.1%, the combined cleaner tailings amounting to 20.4% and the combined rougher tailings amounting to 14.5%, based upon weight of the head sample. The combined cleaner concentrates analyzed 45.67% C1203, the combined cleaner tailings analyzed 19.18% Cr O and the combined rougher tailings analyzed 13.17% Cr O Thus, the sample was up-graded from 36.91% to 45.67% Cr O and 83.63% of the total Cr O was present in the combined concentrates, with 16.37% in the combined cleaner and rougher tailings.

Another sample of the same ore (lot 3240, test 5A- B-C) was treated in the same manner, with the same reagents, in approximately the same amounts, but with additional cleaning of the rougher concentrates of both the sand and slime fractions.

The products from separately floating the sand and slime fractions were combined, with the combined cleaner concentrates amounting to 51.1%, the combined cleaner EXAMPLE XXV To illustrate effectiveness of the process on titanium oxide ores, a sample of ilmenite (Fe, Ti) O was treated. A sample (lot 3313, test 38) assaying 21.1% Ti and which had been grounded to 93.9% 200 mesh, was

floated at 10% solids by weight, with 10 minutes condi;

tioning, minutes rougher flotation, 3 minute first clean ing of concentrate and 3 minute second cleaning of concentrate.

During conditioning, the following reagents were introduced, in the following amounts (pounds per ton of ore): Swift hard soap, 8.9; No. 1 gas oil, 26.6; Oronite S, 0.22; and cuprous sulfate, 0.81. The soap and oil were added as an emulsion. Sodium fluoride, 0.41; and hexametaphosphate, 0.41, were also added, for gangue dispersion and depression.

The concentrate amounted to 43.20%, the cleaner tailings amounted to 33.10% and the rougher tailing amounted to 23.70%, based upon weight of the head sample. The concentrate, cleaner tailings and rougher tailing respectively analyzed 36.40%, 13.84% and 6.95% TiO 71.60% of the total Ti0 was in the concentrate, with 28.40% in the combined cleaner and rougher tailings.

Another sample of the same ilmenite ore (lot 3313, test 41) was treated in the same manner, with the same reagents, in substantially the same amounts, except the cuprous sulfate was omitted and was not replaced by another heavy metal salt.

The concentrate amounted to 11.49%, the cleaner tailings amounted to 29.20% and the rougher tailing amounted to 59.31%, based uponweight of the head sample. The concentrate, cleaner tailings and rougher tailing respectively analyzed 36.20%, 29.40% and 1.300%

14 TiO Only 20.33% of the total TiO; was in the concentrate, with 79.67% in the combined cleaner and rougher tailings.

Thus, when employing a heavy metal salt, the concentrate represented 71.60% of the total TiO but when metal salt was omitted the concentrate represented only 20.33% of the total TiO EXAMPLE XXVI To illustrate the effectiveness of the process on uranium oxide ores, samples of uranium ores from the Colorado Plateau Area near Naturita, Colorado were treated. These ores are so-called'sandstone ores, and the principal uranium mineral identified was carnotite [K (UO )(VO -13H Ol. Carnotite is sometimes classified as a vanadate, but is a uranium oxide ore as termed herein since the uranium value occurs in oxide form. Carnotite is highly friable and slimes easily.

A low-grade sample (lot; 3175, test No. 2) assaying only 0.055% U 0 was ground to 91-93% 200 mesh in a water pulp at 61% solids by weight. The following reagents were introduced during grinding, in the following amounts (pounds per ton of ore): Swift hard soap, 3.75; fuel oil, 14.10; Oronite S, 1.60; and manganous sulfate, 2.00. v

The pulp was floated at 18% solids by weight, for 5 minutes. The concentrate amounted to 17.5% and the tailing amounted to 82.5%, based upon weight of the head sample. The concentrate analyzed 0.271% U 0 and the tailing analyzed 0.019% U 0 Thus, the sample was up-graded from 0.055% to 0.271% U 0 and 75.16% of the total U 0 was present in the concentrate, with 24.84% in the tailing.

The same procedure was repeated, except the sample was a higher grade ore, and a rougher concentrate was cleaned to produce a cleaner concentrate and a cleaner tailing. The sample (lot 3176, test 4-B) assayed 0.236% U 0 The reagents were the same, in substantially the same amounts. Rougher flotation was for 5 minutes, and rougher concentrate was cleaned for 4 minutes.

The cleaner concentrate amounted to 10.1%, the.

cleaner tailing amounted to 13.1% and the rougher tailing amounted to 76.8%, based upon weight of the head sample. The concentrate analyzed 1.156% U 0 the cleaner tailing analyzed 0.389% U 0 and the rougher tailing analyzed 0.08% U 0 Thus, the sample was up-graded from 0.229% to 1.156% U 0 and 50.95% of the total U 0 was present in the concentrate, with 49.05% in the combined cleaner and rougher tailings.

In'this test the rougher concentrate, before cleaning, amounted to 23.2% by weight of the crude ore, and would assay 0.722% U 0 and contain 73.19% of the total U308- The procedure for concentrating uranium of the mineral carnotite [K (UO (VO -13H O] is applicable for concentrating other uranium ores such as uraninite and pitchblende (U0 autunite and uranophane [Ca(UO Si O -6H O]. While antunite and uranophane may be classified as phosphate and silicate minerals, they are oxide ores within the definition of the term as employed herein, in that the uranium occurs in its oxide form, and treatment by the described process up-grades and increases recovery of uranium oxide in the flotation concentrate.

My process is characterized by a high degree of mineral flocculation on all ores I have tried. With some ores, especially those having a high silica or lime content, the gangue also appears to become activated and flocculated. When those conditions prevail, selectivity is poor. I have found particularly in the case of a Mesabi iron washery tailing that an increase in the amount of wetting l agent appeared to increase thedegree of gangue flocculation. Sodium silicate also showed this eifect. However, I have been able to effect gangue deflocculation on such iron pulp when employing my process using sodium fiuosilicate in the conditioning step after making the pulp acid to about pH 4.0.

I have found that in successful operation the fine gangue slimes in the rougher float are highly dispersed. Occasionally this occurs without a wetting or gangue dispersing agent, but in the great majority of cases dispersion is most readily effected by organic wetting agents. In some cases, inorganic gangue dispersing agents such as sodium silicate, sodium fluosilicate, sodium iluoborate, sodium phosphate, sodium carbonate, sodium .borate; or an acid, such as sulfuric, hydrofluoric or fiuosilicic; or a hydroxide, such as sodium hydroxide, either alone or in combination with an organic wetting agent has proven necessary or has producedimproved results. The acids may also function to provide a heavy metal salt, as has been hereinabove. described. 7

In some cases I have noted that increasing the amount of wetting agent increases frothing and at the same time the froth tends to become more brittle, with the result that coarse ore particles tend to drop out into the rougher tailing. On the other hand, if a decreased amount of wetting agent is employed, the rougher froth becomes more voluminous and too stiff with a resulting poor recovery. With each ore to be concentrated the optimum amount of wetting agent will vary and must be determined by preliminary experimentation. I have found that various wetting agents known to the art may be successfully employed in place of the Oronite S used in Example I. The following table illustrates the results obtained by substituting several other such wetting agents.

A proprietary product, being the sodium salt of the sulphonation product of the reaction product of tetraisopropylene and a lower homologue of benzene.

Z A, proprietary product, being d-Z-cthyl hexyl sodium sulfosuceinate.

A proprietary product, being a propylated sodium naphthalene sulionate. a

A proprietary product, being the ethanol amine salt of sulfated coconut fatty alcohol.

1 4% proprietary product, being the sodium salt of a suliated fatty a co 0 QA proprietary product,

being an alkanolamine salt of sulfonated complex alcohols.

Other wetting agents are set forth in the book Surface Active Agents by Schwartz and Perry, published by Interscience Publishers, New York 1949. In addition, the Bureau of Mines Bulletin 449 entitled Development and Use of Certain Flotation Reagents by Deanand Ambrose, Washington 1944, lists othenwetting agents.

A further optional function of the wetting agent is for emulsification of the organic acid collector and/or the hydrocarbon oil, prior to the addition thereof to the pulp, whereby such addition to the pulp is mechanically facilitated.

I have alsofo'und that various other hydrocarbon oils known to be useful in the art of froth flotation may be used in'place of the fuel oil used in the preceding examples; The following table illustrates the results obtained employing several such oils.

. particular ones shown in the examples.

In general, any of the usual hydrocarbon oils used in flotation can be substituted for the particular ones disclosed in the examples. Thus, among the useful hydrocarbon oils are crude petroleum, kerosene, gas oil, domestic fuel oil, coal and wood tars and creosotes.

The practice of the present process is in accordance with the usual procedures in the flotation art, with the exception of the particular combination of reagents used, and hence numerous variations may be made in the particular procedures described in the examples to provide other embodiments which fall within the broad scope of the present process. Thus, in place of the particular ores used in the examples, there may be substituted any of the heavy metallic oxide ores. In preparing the ore for flotation, the ore is first reduced to a powder, as is customary in this art, the powder usually being suificiently fine to pass a mesh sieve. In preparing the pulp, it is not essential that softened water he used, since the present process has been operated using untreated hard water taken from Lake Mead, Nevada, with similar results. Thus, the process of my invention was conducted in a semi-commercial plant in which the general procedure employed in Example I was employed using ore of the same source and analysis, ground to within 1% above 65-rnesh. The pulp was diluted with raw Lake Mead water to an average solids content of 10%.

Since the equipment was inadequate for proper conditioning, an additional conditioning stage was employed prior to the refiotation of the rougher concentrate. The last cleaner tailing (third middling) was recirculated. The following rates and reagents were employed:

Pounds per hour (The manganous sulfateused was an impure product analyzing about 65% MnSO The duration of the test was 7 hours and the following results were obtained:

The final concentrate amounted to 52.0% based on the weight of ore charged and analyzed 41.80% manganese, 9.05% silica and 3.40% alumina. The first and second rniddlings amounted to 7.5% and 5.4%, respectively, and analyzed 13.63% and 17.82% manganese. The tailings amounted to 35.1% based on the weight of the ore charged and analyzed 4.95% manganese.

Thus, the final concentrate represented a recovery of 85.4% of the manganese ore while 6.8%, 4.0% and 3.8% were in the tailings, first middling and second middling, respectively.

The specific examples illustrate useful pulp concentrations, but the particular concentration employed is not critical. In general, the slimier the ore, the more dilute the flotation rougher feed should be.

Any of the usual carboxylic acid or soap collectors such as fatty acid collectors containing at least 12 carbon atoms and naphthenic acids may be substituted for the As the collector, there can be used any of the long chain carboxylic acids,

1 amples.

such as lauric, palmitic, stearic, oleic or ricinoleic acid, or mixtures containing such acids ,for example, mixtures obtained by the splitting of cotton seed oil, coconut oil, palm oil, castor oil, and so forth. The ammonium, alkali metal and amine (e. g., ethyl amine) soaps (i. e. salts) of such acids or of the mixtures containing such acids are also useful collectors.

.Any heavy metal salt which is water soluble can be substituted for the particular ones disclosed in the ex- Thus, salts of cobalt, nickel or zinc have also been used. All of the heavy metal salts have not been found to be equally effective with all ores. With manganese ores it was found preferable to use a salt of manganese, copper, lead or iron.

The relative amounts of collector, heavy metal salt, oil and wetting agent used are not critical and can be varied, but with some effect on the effectiveness of the process. cess of the salt, based upon the amount of fatty acid collector present.

The oxide ore of a heavy metal, to which the invention is applicable and as this expression is used in the claims, may be a mineral containing an oxide of the heavy metal, e. g. wad, which is a mixture of hydrous manganese oxides, expressed in some mineral tables as MnO H O, from which manganese is the value to be recovered, or uraninite, U from which uranium is the value to be recovered; or an oxide of the heavy metal and one or more other elements, e. g. wolframite (Fe, Mn) W0 from which tungsten is the value to be recovered, or ilmenite (Fe, Ti) O from which titanium is the value to be recovered; or an oxide of the heavy metal together with an oxide of one or more other elements, e. g. carnotite [K (UO (VO -13H O], or autunite [Ca(UO (PO -1O12H O], or uranophane from which uranium is the value to be recovered.

While, for purposes of mineral classification, Wolframite may be listed as a tungstate, carnotite may be listed as a vanadate, autunite may be listed as a phosphate and uranophane may be listed as a silicate, they are oxide ores within the meaning of the invention, in that the heavy metal value occurs in oxide form, whether or not it is a simple oxide, e. g. MnO U0 etc., or a complex oxide, e. g. (Fe, Ti) O (Fe, Mn) W0 [K2(UO2)2(VO4)2' etc. I

A number of additional uranium minerals are listed in U. S. Geological Survey Bulletin 1009-F, entitled Glossary of Uraniumand Thorium-Bearing Minerals, Ed. 3, Washington 1955. These are substantially all simple or complex oxide ores of uranium, with or without accessory metals, and are thus heavy metal oxide ores within the meaning of that term as used herein.

This application is a continuation-in-part of my copending application, Serial No. 149,447, filed March 13, 1950, now abandoned, and my copending application Serial No. 215,645, filed March 14, 1951, now abandoned.

I claim:

1. In the concentrating of a manganese oxide ore by flotation, the step of incorporating into the pulp an organic acid collector selected from the group consisting of fatty acids having at least 12 carbon atoms and naphthenic 'acids and their corresponding water soluble soaps, a water soluble salt of a heavy metal, a hydrocarbon oil, and a wetting agent.

2. A process as in claim 1 in which the heavy metal salt is a manganese salt.

3. A process as in claim 1 in which the heavy metal salt is selected from the group consisting of iron, cobalt and nickel salts.

4. A process as in claim 1 in which the heavy metal salt is an iron salt.

5. A process as in claim 1 in which the heavy metal salt is a copper salt.

It is also preferred to use a stoichiometric ex- 6. A process as in claim 1 in which the heavy metal salt is a lead salt.

7. The process as in claim 1 in which the collector is sodium oleate.

8. The process as in claim 1 in which the wetting agent is an alkyl aryl sulfonate.

9. The process as in claim 1' in which the hydrocarbon oil is a petroleum oil.

10. The process as in claim 1 in which the heavy metal salt is a manganese salt and in which the collector is sodium oleate.

11. The process as in claim 1 in which the heavy metal salt is selected from the group consisting of iron, cobalt and nickel salts and in which the collector is sodium oleate.

12. The process as in claim 1 in which the heavy metal salt is a copper salt and in which the collectoris sodium oleate.

13. The process as in claim 1 in which the heavy metal salt. is a lead salt and in which the collector is sodium oleate.

14. The process as in claim 1 in which the heavy metal salt is a manganese salt, in which the collector is sodium oleate and in which the wetting agent is an alkyl aryl sulfonate.

15. The process as in claim 1 in which the heavy metal salt is selected from the group consisting of iron, copper and nickel salts, in which the collector is sodium oleate and in which the wetting agent is an alkyl aryl sulfonate.

16. The process as in claim 1 in which the heavy metal salt is a copper salt, in which the collector is sodium oleate and in which the wetting agent is an alkyl aryl sulfonate.

17. The process as in claim 1 in which the heavy metal salt is a lead salt, in which the collector is sodium oleate and in which the wetting agent is an alkyl aryl sulfonate.

18. The process as in claim 1 in which the heavy metal salt is a manganese salt, in which the collector is sodium oleate, in which the Wetting agent is an alkyl aryl sulfonate and in which the hydrocarbon oil is a petroleum oil.

19. The process as in claim 1 in which the heavy metal salt is selected from the group consisting of iron, cobalt and nickel salts, in which the collector is sodium oleate, in which the wetting agent is an alkyl aryl sulfonate and in which the hydrocarbon oil is a petroleum oil.

20. The process as in claim 1 in which the heavy metal salt is a copper salt, in which the collector is sodium oleate, in which the Wetting agent is an alkyl aryl sulfonate and in which the hydrocarbon oil is a petroleum oil.

21. The process as in claim 1 in which the heavy metal salt is a lead salt, in which the collector is sodium oleate, in which the wetting agent is an alkyl aryl sulfonate and in which the hydrocarbon oil is a petroleum oil.

22. The process as in claim 1 in which the heavy metal salt is an iron salt, in which the collector is sodium oleate, in which the wetting agent is an alkyl aryl sulfonate and in which the hydrocarbon oil is a petroleum oil.

23. The process as in claim 1 in which the wetting agent is a fatty alcohol sulfate.

24. The process as in claim 1 in which the manganese oxide ore is a slimy ore, in which the heavy metal salt is a manganese salt, in which the collector is sodium oleate, in which the wetting agent is an alkyl aryl sulfonate and in which the hydrocarbon oil is a petroleum oil.

25. The process as in claim 1 in which manganese oxide ore is a slimy ore, in which the heavy metal salt is a manganese salt, in which the collector is sodium oleate, in which the wetting agent is an alkyl aryl sulfonate, in which the hydrocarbon oil is a petroleum oil, and in which the collector and wetting agent are emulsified together before being incorporated into the pulp.

26. In concentrating by a flotation process an ore which is an oxide of a heavy metal of the group consisting of titanium, iron, manganese, chromium, tin and uranium, the step of incorporating into the pulp an organic acid collector selected from the group consisting of fatty acids having at least 12 carbon atoms and naphthenic acids and their corresponding water soluble soaps, a water soluble salt of a heavy metal, a hydrocarbon oil, and a wetting agent.

27. In the concentration of a manganese oxide ore by flotation, the step of incorporating into the pulp an organic acid collector selected from the group consisting of fatty acids having at least 12 carbon atoms and naphthenic acids and their corresponding water soluble soaps, a water soluble salt of a heavy metal and a hydrocarbon oil.

28. In concentrating by a flotation process an ore which is an oxide ore of a heavy metal of the group consisting of titanium, iron, manganese, chromium, tin and uranium, wherein said ore is formed into a pulp and subjected to a flotation step, the step of incorporating into the pulp an organic acid collector selected from the group consisting of fatty acids having at least 12 carbon atoms and naphthenic acids and their corresponding water References Cited in the file of this patent UNITED STATES PATENTS 1,911,865 Weinig May 30, 1933 1,951,326 De Vaney Mar. 13, 1934 1,955,039 Weinig et a1. Apr. 17, 1934 1,979,324 Gaudin Nov. 6, 1934 2,120,485 Clemmer et al. June 14, 1938 2,297,664 Tartaron et a1. Sept. 29, 1942 2,607,479 Bates Aug. 19, 1952 OTHER REFERENCES Flotation Fundamentals, part I, by A. N. Gaudin et a1., published in 1928 by the University of Utah, page 97.

Taggart: Handbook of Mineral Dressing, 1945, John Wiley and Sons, Inc., New York, pages l213, 12109, 12-113, 12115, 12119, 12-120. 

28. IN CONCENTRATING BY A FLOTATION PROCESS AN ORE WHICH IS AN OXIDE ORE OF A HEAVY METAL OF THE GROUP CONSISTING OF TITANIUM, IRON, MANGANESE, CHROMIUM, TIN AND URANIUM, WHEREIN SAID ORE IS FORMED INTO A PULP AND SUBJECTED TO A FLOTATION STEP, THE STEP OF INCORPORATING INTO THE PULP AN ORGANIC ACID COLLECTOR SELECTED FROM THE GROUP CONSISTING OF FATTY ACIDS HAVING AT LEAST 12 CARBON ATOMS AND NAPHTHENIC ACIDS AND THEIR CORRESPONDING WATER SOLUBLE SALTS, A WATER SOLUBLE SALT OF A HEAVY METAL AND A HYDROCARBON OIL. 